Process for the extraction of zinc from sulphide concentrates

ABSTRACT

A portion of the concentrates is roasted and a portion of the resulting calcine is subjected to a neutral leaching step. Another portion of the concentrates is directly leached in an acidic medium in the presence of oxygen and under atmospheric conditions together with at least a portion of the neutral leaching residue. The zinc- and iron-rich solution resulting from acid leaching is neutralised with another portion of the calcine, the iron is removed and the solution is reused in the neutral leaching step. The method enables a gradual increase in the capacity of an existing zinc plant in accordance with demand, while capital costs may advantageously be spread out over time.

FIELD OF THE INVENTION

The present invention relates to a process for the extraction of zincfrom zinc sulphide concentrates, comprising the following operations:

(a) roasting a portion of the concentrates so as to produce calcine,

(b) neutral leaching of calcine produced in the operation (a) withreturn spent electrolyte so as to produce a leachate which is rich inzinc and substantially iron-free, which is separated, and a zinc ferriteresidue, which is separated,

(c) leaching of another portion of the concentrates and of at least aportion of the zinc ferrite residue produced in the operation (b) with asolution of sulphuric acid returning from electrolysis at 60°-95° C. inatmospheric conditions and in the presence of finely dispersed oxygen soas to produce a leachate which is rich in zinc and in iron, which isseparated, and a leaching residue which is depleted in zinc, which isseparated, the quantities of concentrates and of zinc ferrite which areused in this operation (c) being such that the molar ratio of the ironcontained in the ferrite and the reactive sulphur contained in theconcentrates is at least approximately 0.2,

(d) conditioning, preparatory to the following operation, of theleachate which is rich in zinc and in iron, produced in the operation(c),

(e) precipitation of the major portion of the iron contained in thesolution conditioned in the operation (d) so as to produce a solutionwhich is rich in zinc and depleted in iron and a ferriferousprecipitate, which is separated, and

(f) introduction of the solution which is rich in zinc and depleted iniron into the neutral leaching in (b).

The following should be understood here by

"zinc sulphide concentrate": a concentrate containing, in the form ofsulphides, chiefly zinc and iron and, in smaller proportions, copper,silver and/or lead;

"in atmospheric conditions": in conditions which do not require the useof autoclaves, that is to say at a pressure equal to or differing byless than 20 kPa from atmospheric pressure; and

"reactive sulphur": the sulphur present in the form of sulphide in thezinc sulphide concentrates and in the leaching residue which is rich inzinc (an expression employed later) and which can be oxidized by ferricsulphate according to the reaction:

    Fe.sub.2 (SO.sub.4).sub.3 +MeS=MeSO.sub.4 +2FeSO.sub.4 +S°(I)

in which Me denotes Zn, Fe, Cu, Pb or Ag. (the reactive sulphurgenerally consists of all the sulphur present in the form of sulphideless the pyrite sulphur).

Using the process of the invention, a leachate which is rich in zinc andsubstantially iron-free, a leaching residue which is depleted in zincand a ferriferous precipitate are therefore obtained. This leachate canbe purified and subsequently electrolysed in order to extract the zincfrom it. The leaching residue which is depleted in zinc, and whichcontains sulphur, lead sulphates, silver compounds, undissolvedsulphides (pyrites) and gangue, can be subjected to an appropriatetreatment in order to extract the sulphur and the valuable metals fromit. The ferriferous precipitate can be stored or, when it is pureenough, can be employed as pigment or as source of iron in the steelindustry.

BACKGROUND

A process similar to that as defined above is described in documentEP-A-0451456. In this known process, all of the calcine produced in theoperation (a) is leached in the operation (b) and all of the ferriteproduced in the operation (b) is leached in the operation (c), whileemploying a concentrate:ferrite ratio such that approximately 15 to 20%of the trivalent iron, required for the oxidation according to reaction(I) of the reactive sulphur present in the concentrate, originates fromthe leaching of the ferrite according to the reaction

    ZnO.Fe.sub.2 O.sub.3 +4H.sub.2 SO.sub.4 =Fe.sub.2 (SO.sub.4).sub.3 +ZnSO.sub.4 +4H.sub.2 O                                   (II)

The remainder of the trivalent iron required for the oxidation of thereactive sulphur is obtained by the reaction

    2FeSO.sub.4 +H.sub.2 SO.sub.4 +0.50.sub.2 =Fe.sub.2 (SO.sub.4).sub.3 +H.sub.2 O                                                (III)

It is proposed to work in (c) in such a manner that the leachate whichis rich in zinc and in iron has a sulphuric acid content of 10-25 g/land an Fe³⁺ content of less than 10 g/l, which is apparentlyunobtainable in a single leaching stage. This is why the leaching in (c)is carried out in two stages.

In the first stage, the zinc ferrite and a leaching residue which isrich in zinc, produced in the second stage, are treated with thesolution of acid returning from electrolysis so as to produce a primaryleachate containing 50-90 g/l of H₂ SO₄ and the leaching residue whichis depleted in zinc, which are separated. No oxygen is employed in thisfirst leaching stage, this being to make it possible to use in thisstage simpler types of reactors than in the second stage. Theinvolvement of reaction (III) is therefore not brought about in thefirst stage.

In the second stage, the concentrates are treated with the said primaryleachate in the presence of finely dispersed oxygen so as to produce, byreactions (I) and (III), a leachate which is rich in zinc and in ironand a leaching residue which is rich in zinc. At the end of this secondstage, the operation (d) is performed by adding a small quantity offresh concentrate to the leaching pulp so as to convert ferric sulphateinto ferrous sulphate by reaction (I); the operation (d) is thereforeincorporated into the operation (c) and, as a result of the secondleaching stage, there is obtained the leaching residue which is rich inzinc, which is separated and recycled into the first stage, and aleachate which is rich in zinc and in iron, which is alreadyconditioned.

The operation (e) is performed by adding more concentrate to theconditioned solution and by then precipitating the iron in haematiteform by oxidation in an autoclave.

This produces, on the one hand, the said solution which is rich in zincand depleted in iron and, on the other hand, a precipitate of haematitecontaining a small quantity of elemental sulphur and of sulphides. Thissulphur and these sulphides are subsequently separated from thehaematite by flotation. The reason why the concentrate is used in theoperation (e) is not given. A possible explanation could be that theacidity of the conditioned solution is too high to permit a suitableprecipitation of the iron and that, because of this, concentrate isadded as neutralizing agent (reactions (I) and (II)).

This known process therefore requires leaching in two stages and, sincethe work is carried out with a concentrates:ferrite ratio such thatapproximately 15 to 20% of the trivalent iron, required for theoxidation according to reaction (I) of the reactive sulphur present inthe concentrate, originates from the leaching of the ferrite inaccordance with reaction (II) and that oxygen is not employed in thefirst leaching stage, it is necessary to oxidize approximately 80 to 85%of the reactive sulphur in the second leaching stage using trivalentiron obtained by reaction (III), when a zinc leaching yield close on100% is aimed at.

However, the Applicant has found that in these conditions the secondleaching stage takes place very slowly, and this obviously constitutes aserious disadvantage. Another disadvantage of this known process lies inthe fact that the operation (c) is not easy to control because theleaching yield is determined solely by the ratio of the reactive sulphurto the zinc ferrite which are introduced into the first leaching stageand because the system reacts very slowly to corrections which are madeto this ratio.

Furthermore, the use of this known process in existinghydrometallurgical zinc plants would almost always entail a considerableinvestment for purchasing the autoclaves required for the operation (e).In fact, to the Applicant's knowledge, there are only two plants in theworld which make haematite and which are therefore already equipped withsuch autoclaves; all the others make jarosite or goethite in atmosphericconditions and are therefore not endowed with such autoclaves.

Moreover, as in this known process all of the ferrite produced in theoperation (b) is leached in (c) while employing a concentrate:ferriteratio such that approximately 15 to 20% of the trivalent iron, requiredfor the oxidation of the reactive sulphur, originates from the leachingof the ferrite, the installation of this process in an existing plant,the roasting capacity of which would quite logically be maintained atthe existing level, would have to result at once in approximatelydoubling the plant capacity. However, an increase in the capacity of anexisting plant which is as substantial as this all at once, which willbe unavoidably accompanied with substantial investments, will not oftenbe opportune. The process therefore lacks some degree of flexibility.

What is more, this known process produces a haematite which is soiledwith sulphur and sulphides.

SUMMARY OF THE INVENTION

The objective of the present invention is to provide a process asdefined above which avoids the disadvantages of the known process.

To this end, according to the invention

(1) only a portion of the calcine produced in the operation (a) isleached in the operation (b),

(2) the leaching in (c) is performed

either in a single stage, in which case the work is done so that theleachate which is rich in zinc and in iron has a sulphuric acid contentof 45-75 g/l, preferably of 55-65 g/l, and an Fe³⁺ content of 1-10 g/l,preferably of 2-5 g/l,

or in two stages, in which case the work is done so that the leachatewhich is rich in zinc and in iron has a sulphuric acid content of 10-35g/l, preferably of 10-25 g/l, and an Fe³⁺ content of 0.1-2 g/l,preferably of 0.5-1 g/l, the first stage comprising treating the zincferrite and a leaching residue which is rich in zinc, produced in thesecond stage, with the solution of sulphuric acid returning fromelectrolysis in the presence of finely dispersed oxygen so as to producea primary leachate and the said leaching residue which is depleted inzinc, which are separated, and the second stage comprising treating theconcentrates with the said primary leachate in the presence of finelydispersed oxygen so as to produce the said leachate which is rich inzinc and in iron and the said leaching residue which is rich in zinc,which are separated, this second stage being performed in conditionssuch that less than 60% and preferably less than 50% of the reactivesulphur are oxidized therein and that the leaching residue which is richin zinc produced therein has a reactive sulphur content which isappreciably higher than that which can be oxidized in the first stage bythe iron present in the ferrite,

(3) the operation (d) is performed by treating the leachate which isrich in zinc and in iron with yet another portion of the concentrates soas to return its Fe³⁺ content below 5 g/l and preferably to 1-3 g/l, itbeing possible for this reducing treatment to be omitted when theleachate which is rich in zinc and in iron already has an Fe³⁺ contentof less than 5 g/l, and by treating the solution of low Fe³⁺ contentwith another portion of the calcine produced in the operation (a), so asto return the free H₂ SO₄ content of this solution below 10 g/l andpreferably to 3-5 g/l, this neutralization treatment producing, on theone hand, a zinc ferrite residue, which is separated and subsequentlytreated in the same way as the ferrite produced in the operation (b)and, on the other hand, a conditioned solution,

(4) the operation (e) is performed by precipitating the iron in a mannerwhich is known per se in the form of goethite, haematite, jarosite orother compound which has suitable filterability, this being in theabsence of zinc sulphide concentrate, and

(5) the leaching in the operation (c) is carried out on

either all of the ferrite residue produced in the operations (b) and(d),

or only a portion of this ferrite, in which case the remainder of thisferrite is treated separately in a manner which is known per se by hotacidic leaching, this treatment producing another leachate which is richin zinc and in iron, and this solution is subjected to the operations(d), (e) and (f) together with the leachate which is rich in zinc and iniron, produced in the operation (c).

In fact, by leaching in the operation (b) only a portion of the calcineproduced in the operation (a), it is possible to employ another portionof this calcine in the operation (d).

By performing the operation (c) as defined in (2), the leaching periodis appreciably shortened, whether working in only one or in two stages,as will be demonstrated later, and this operation (c) can be easilycontrolled, given that the leaching yield is now determined by thequantity of oxygen used and that the system reacts promptly tocorrections which are made to this parameter.

By performing the operation (d) as defined in (3), a conditionedsolution is obtained in which the iron can be precipitated by anyconventional oxidation and hydrolysis technique, this being done with aminimum of neutralizing agent, when goethite or jarosite isprecipitated, and without it being necessary to add zinc sulphideconcentrate, when haematite is precipitated.

By performing the operation (e) as defined in (4), the process of theinvention can be installed in any existing hydrometallurgical zinc plantwhatever, without this necessarily entailing a large investment.

Owing to its characteristic defined in (5), the process of the inventionmakes it possible to increase gradually the capacity of an existingplant, this being according to the needs and with an advantageousstaging of the investment costs.

It is appropriate that what follows should be reported here.

Documents U.S. Pat. No. 3,976,743, U.S. Pat. No. 4,107,265 andBE-A-724214 describe processes for the treatment of zinc ferrite whichmake use of reactions (I) and (II), but not of reaction (III). Theseknown processes do not make it possible to increase the capacity of theexisting zinc plants producing ferrite, because all these plants alreadyutilize reactions (I) and (II) in one way or another.

Document WO-A-91/09146 describes a process for the treatment of zincferrite, comprising, successively, leaching of the ferrite with acidreturning from electrolysis (reaction II), partial neutralization of theresidual acid by addition of ZnS concentrate in the presence of oxygen(reactions I and III), reduction of the trivalent iron by addition ofconcentrate (reaction I), flotation of the pulp so as to separate fromit elemental sulphur and unreacted concentrate, treatment of theflotation residue with SO₂ in order further to leach the iron, the zincand the impurities, treatment of the pulp resulting therefrom withelemental sulphur to precipitate the copper, flotation of the pulp so asto separate from it a copper sulphide concentrate, filtration of thepulp and precipitation of the iron in the resultant solution. This knownprocess differs from the process of the invention not only in itscomplexity but also in the fact that reaction (II) is used beforereactions (I) and (III), which lengthens the leaching period, as theApplicant has ascertained.

Documents U.S. Pat. No. 4,510,028 and EP-A-0071684 describe a processfor the treatment of zinc ferrite by acidic leaching in one or twostages, in the presence of concentrate and with oxygen under pressure at135°-175° C. (reactions I, II and III). The ferrite:concentrate ratiomust be such that the zinc contained in the ferrite amounts to 5-40% andpreferably to 5-20% of all the zinc contained in the ferrite and theconcentrate. In contrast to the process of the invention, this knownprocess therefore requires autoclaves for leaching the ferrite and theconcentrate. Moreover, since this known process gives the best resultswith a low ferrite:concentrate ratio, its installation into an existingplant producing ferrite would at once enormously increase the capacityof this plant, which is not often opportune.

Document EP-A-0166710 describes a process as defined at the beginning ofthe present application, except that the concentrates:ferrite ratioemployed in the operation (c) is not specified, that the operation (c)is performed partially under pressure and that the operation (d) isomitted. In this known process, a portion of the calcine produced in theoperation (a) is leached in the operation (b) and all of the ferriteproduced in the operation (b) is leached in the operation (c). Theoperation (c) is performed in three stages. In the first stage, theferrite and a leaching residue which is relatively depleted in zinc,produced in the second stage, are treated with acid returning fromelectrolysis in the presence of oxygen and in atmospheric conditions soas to produce a primary leachate and a leaching residue which isdepleted in zinc, which are separated. In the second stage, a leachingresidue which is rich in zinc, produced in the first stage, andoptionally concentrate, are treated with the said primary leachate inthe presence of oxygen and at 120°-160° C., that is to say in anautoclave or equivalent apparatus, so as to produce a secondary leachateand the said residue which is relatively depleted in zinc, which areseparated. In the third stage, concentrate is treated with the saidsecondary leachate in the presence of oxygen and in atmosphericconditions so as to produce a leachate which is rich in zinc and in ironand the said leaching residue which is rich in zinc, which areseparated. The work is done so that the leachate which is rich in zincand in iron has an acid content of approximately 4 to 8 g/l. Thissolution is subjected directly to the operation (e), which consists inprecipitating the iron in the form of goethite, using as neutralizingagent the other portion of the calcine produced in (a). This knownprocess differs from the process of the invention not only in theabsence of the operation (d) and the complexity of the operation (c),the use of which additionally requires an autoclave or equivalentapparatus, but also in the fact that virtually all of the acid isexhausted in the operation (c) by the reactions (I) and (III). However,it has been found that the overall duration of the operation (c) is thuslengthened excessively. Moreover, as in this known process the otherportion of the calcine produced in (a) is employed as neutralizing agentin (e), goethite containing a substantial quantity of zinc ferrite isnecessarily produced, and this can be avoided in the process of theinvention.

Document U.S. Pat. No. 4,004,991 describes a process for the extractionof zinc from sulphide concentrates, according to which the concentratesare leached in two stages countercurrentwise with acid returning fromelectrolysis in the presence of oxygen at 135°-175° C., that is to sayin an autoclave. As this known process does not comprise the operations(a) and (b), the only point in common between this process and theprocess of the invention lies in the fact that a leaching is performedin two stages with acid returning from electrolysis.

When the operation (e) is excluded, the process of the inventionprovides for four different routes, which will be called "variants"below:

first variant:

performing the operation (c) in a single stage with only a portion ofthe ferrite residue produced in the operations (b) and (d)

second variant:

performing the operation (c) in a single stage with all of the ferriteresidue produced in the operations (b) and (d)

third variant:

performing the operation (c) in two stages with only a portion of theferrite

fourth variant:

performing the operation (c) in two stages with all of the ferrite.

When working in comparable conditions (the same concentrate and the samequantity of concentrate employed in (a), the same molar ratio of iron inthe ferrite to the reactive sulphur in the concentrate employed in (c)and, when the first and the third variants are employed, the samefraction of ferrite used in (c)), the zinc output will be the lowest inthe first variant and the highest in the fourth. In the first and thirdvariants, the zinc output can be varied with the fraction of ferriteused in (c). In each of the four variants, the zinc output can also bevaried by modifying the said molar ratio. As already mentioned above,the conventional process for the extraction of zinc, employed in theexisting plants which make ferrite, already makes use of the reactions(I) and (II). The increase in output which is obtained by substitutingthe process of the invention for this conventional process in theseplants will therefore be linked essentially with the quantity of zincdissolved in (c) by the reactions (I) and (III). The first variant willtherefore be employed when it is intended to produce a relatively smallincrease in capacity (for example from 5 to 10%) or when it is intendedto produce a number of increases of small extent consecutively. Thesecond variant will be employed to increase the plant capacitysubstantially, and the fourth when it is intended to increase thecapacity further. The third variant will normally be employed only whenit is intended, for any reason whatever, to continue to treat a portionof the ferrite by the conventional route and at the same time to drawmaximum profit from the fraction of ferrite used in (c).

The molar ratio of the iron contained in the zinc ferrite to thereactive sulphur contained in the concentrate is at least approximately0.2 and preferably at least 0.3 in order that the rate of leaching in(c) should not become too low. It is obvious that this ratio must belower than 2 in order that it may still be possible to resort to thereaction (III). In the fourth variant, this ratio will be advantageouslyequal to or lower than 0.6, preferably equal to or lower than 0.4, inorder that the zinc output should be at a maximum. This ratio of ≦0.6 isfurthermore also suitable in the case of the other variants.

With regard to the conditions of the leaching in one stage (first andsecond variants):

the H₂ SO₄ content of the leachate which is rich in zinc and in iron isat least 45 g/l and preferably at least 55 g/l; otherwise, there is arisk of precipitating lead and silver jarosites which not only interferewith the leaching itself but can moreover subsequently be detrimental tothe recovery of precious metals from the zinc residue; furthermore, anacid content which is too low also complicates the separation of theresidue which is depleted in zinc from the leachate;

the H₂ SO₄ content of the leachate is not higher than 75 g/l andpreferably not higher than 65 g/l; otherwise too much calcine must beemployed in (d);

the Fe³⁺ content of the leachate is 1-10 g/l, preferably 2-5 g/l,because in these conditions the leaching rate and yield are optimal.

It is particularly useful to take care that the trivalent ironconcentration does not drop below approximately 0.1 g/l, preferably notbelow 0.2 g/l, during the initial phase of the leaching. If there is adrop below approximately 0.1 g/l of Fe³⁺, there is a risk not only ofhaving corrosion problems, especially with the steels commonly employedfor the construction of leaching equipment, but also of forming H₂ S andof seeing the copper disappear from the solution, copper which catalysesthe reaction III. To avoid these problems, the potential of the pulpmust be at least 530 mV (SHE) and preferably at least 560 mV.Furthermore, it is also advantageous to watch that the potential of thepulp does not rise above 640 mV in the said initial phase, becauseferrite dissolves less quickly above 610 mV.

It is therefore important to check rigorously, especially usingpotential measurements, the trivalent iron concentration of the solutionin the various phases of the leaching and to adjust this concentrationas necessary by modifying the flow rate of oxygen and/or thetemperature, a reduction in the temperature making the reactive sulphurless reducing and therefore less demanding for trivalent iron.

With regard to the conditions of leaching in two stages (third andfourth variants):

the H₂ SO₄ content of the leachate which is rich in zinc and in iron isat least 10 g/l; otherwise the leaching period is appreciablylengthened;

the H₂ SO₄ content of the said leachate is not higher than 35 g/l,preferably not higher than 25 g/l; otherwise too much calcine must beemployed in (d);

the Fe³⁺ content of the said leachate is 0.1-2 g/l, preferably 0.5-1g/l; if there is a drop below 0.1 g/l of Fe³⁺, there is a risk of havingthe abovementioned problems; on the other hand, if there is a rise above2 g/l of Fe³⁺, there is a risk of forming lead and silver jarosites, andthis makes the separation of the leaching residue which is rich in zincfrom the leachate which is rich in zinc and in iron much more difficult.

It is advantageous to oxidize at least 30%, preferably at least 40%, ofthe reactive sulphur in the second stage of leaching. If less than 30%of this sulphur is oxidized in the second stage, there is a risk ofconsuming too much acid in the first leaching stage and thus forminglead and silver jarosites, which not only interfere with the leachingitself but which can furthermore subsequently be detrimental to therecovery of the valuable metals from the leaching residue which isdepleted in zinc.

It is particularly useful to perform the first stage of leaching so thatthe trivalent iron concentration of the solution, which will necessarilybe low during the initial phase of this stage, reaches a value of 2-10g/l, preferably of 3-7 g/l, in the final phase of this stage. It is, infact, in these conditions that the leaching rate and yield becomeoptimal.

It is furthermore important to take care that the trivalent ironconcentration does not drop below 0.1 g/l, preferably not below 0.2 g/l,during the said initial phase, because otherwise there is a risk ofhaving the abovementioned problems: corrosion, formation of H₂ S anddisappearance of the copper from the solution. It is therefore importantto watch that the potential is at least 530 mV and preferably at least560 mV in the said initial phase and it is also important to controlrigorously, especially by potential measurements, the trivalent ironconcentration of the solution in the other phases of the first stage ofleaching and to adjust this concentration as required, as mentionedabove.

As already stated above, it is not advisable to consume too much acid inthe first stage of leaching. In fact, it is appropriate to end thisstage at an acid concentration of 40-70 g/l, preferably of 55-65 g/l. Itis therefore important to watch that the quantities of acid, of ferriteand of sulphur (in the form of leaching residue which is rich in zinc)which are introduced in the first stage of leaching are such that theprimary leachate has a sulphuric acid content of 55-65 g/l. The secondstage of leaching is advantageously performed so as to maintain thetrivalent iron concentration of the solution constantly at the abovelevel of 0.1-2 g/l, preferably of 0.5-1 g/l, this being in order toavoid the abovementioned problems.

Other details and special features of the invention will emerge from thedescription of two embodiments of the process of the invention, which isgiven by way of nonlimiting example and with reference to the drawingsenclosed herewith.

BRIEF DESCRIPTION OF THE DRAWINGS

In these drawings

FIG. 1 shows a diagram of an existing zinc plant employing theconventional process for zinc extraction;

FIG. 2 shows a diagram of an existing zinc plant which has been adaptedfor using an embodiment of the first variant of the process of theinvention;

FIG. 3 shows a diagram of an existing zinc plant which has been adaptedfor using an embodiment of the fourth variant of the process of theinvention;

FIG. 4 shows diagrammatically the plant used for performing theoperations (c) and (d) in the embodiment of FIG. 3; and

FIG. 5 shows, on larger scale and in more detail, a tank of the plant ofFIG. 4.

In these figures, the same reference numbers indicate identicalcomponents.

DESCRIPTION OF THE PREFERRED EMBODIMENTS

The plant shown in FIG. 1 receives a zinc sulphide concentrate 1 asfeed. A portion 1a of this concentrate is roasted in 2 and a portion 3aof the calcine thus produced is subjected in 4 to a neutral leachingwith sulphuric acid returning from electrolysis. The solution 5 leaving4, which is rich in zinc and in iron-free substance, is purified in 6and electrolysed in 7. The residue 8 from the neutral leaching, which iscomposed essentially of zinc ferrite and of gangue, is introduced intothe first stage 9 of a hot acidic leaching in which stage the ferrite isleached with an acidic solution 12 produced in the second stage 10 ofthis hot acidic leaching. In the second stage 10, the residue 11produced in 9 is leached in a very acidic medium with acid returningfrom electrolysis. The residue produced in 10 contains the gangue andinsoluble compounds, especially lead sulphate. The solution 13 producedin 9 is a leachate which is rich in zinc and in iron: approximately 100g/l Zn, 25-30 g/l Fe³⁺ and 50-60 g/l H₂ SO₄. This solution is treated ina reduction stage 14 with a second portion 1b of the concentrate toreturn its Fe³⁺ content below 5 g/l. The residue 15 produced in 14 isrecycled in 2 and the solution 16 of low Fe³⁺ content, produced in 14,is treated in a neutralization stage 17 with a second portion 3b of thecalcine produced in 2 to return its acid content below 10 g/l. Theferrite residue 18 produced in 17 is recycled at 9 and the conditionedsolution 19 produced in 17 is treated in 20 in order to separate most ofthe iron from it, for example in the form of goethite 21. In this case,oxygen is injected in 20 into the solution while the latter is beingneutralized, preferably with pure calcine 22 obtained by roasting pureZnS concentrates, so as not to lose zinc in ferrite form. The solution23 produced in 20, which is a solution rich in zinc and depleted iniron, is recycled at 4.

It has already been proposed in the literature to modify theconventional process described above in the sense that the reductionstage 14 is eliminated and that the second portion 2a of the concentrateis introduced into the first stage 9 of the hot acidic leaching, whichthen becomes a hot reducing acidic leaching.

FIG. 2 shows the plant of FIG. 1 after its adaptation for using thefirst variant of the process of the invention. An additional quantity 1cof the concentrate and a portion 8a of the ferrite are now leached inone stage with the acid returning from electrolysis in the presence ofoxygen at 24 (operation (c)). The remainder 8b of the ferrite is treatedin 9. The leaching residue which is depleted in zinc 25, produced in 24,is treated in 26 in order to extract from it the elemental sulphur S°and the valuable metals 27. When the solution which is rich in zinc andin iron 28, produced in 24, requires a reduction (solid line) it isadded to the solution 13 (or to the hot reducing leaching, when thelatter is present); otherwise it is added to the solution 16 (dottedline).

FIG. 3 shows the plant of FIG. 1 after its adaptation for using thefourth variant of the process of the invention. Since all of the ferrite8 is now treated in the operation (c) and since in the embodiment whichis to be described the operation (d) is incorporated in the operation(c), stages 9, 10, 14 and 17 are eliminated. The operation (c) isperformed in two stages 29 and 30. In the first stage 29, the ferrite 8and the leaching residue which is rich in zinc 31, produced in thesecond stage 30, are leached with returning acid in the presence ofoxygen. The leaching residue which is depleted in zinc 25, produced in29, is treated, as in the plant of FIG. 2, in 26 in order to extract theelemental sulphur S° and the valuable metals. In the second stage 30, anadditional (substantial) quantity 1b of concentrate is leached in thepresence of oxygen with the solution 32 produced in 29. At the end ofthe leaching in 30, a portion 3b of the calcine is added to the pulp soas to bring the acid content of the solution to below 10 g/l, afterwhich the residue 31 is sent to the first stage 29 and the solution 19,which is already conditioned, to stage 20.

It is obvious that the equipment which is released by eliminating stages9, 10, 14 and 17 can, for the most part, be reemployed for making use ofstages 29 and 30.

The plant shown in FIG. 4 comprises a first series of four leachingtanks 33a, 33b, 33c and 33d which are placed in cascade and followed bya solid-liquid separator 34 and a second series of three leaching tanks35a, 35b and 35c, also placed in cascade and followed by aneutralization tank 35d and a solidliquid separator 36. Each tankoverflows into the following tank, except for the tanks 33d and 35dwhich overflow into the separators 34 and 36 respectively. The separator34 comprises a thickener and a filtration apparatus, and the separator36 a filtration apparatus.

The leaching tanks are closed and equipped, as shown in FIG. 5, with afeed inlet 37, an oxygen inlet 38, a spillway 39 and a self-suckingstirrer 40, for example a self-sucking stirrer with a hollow shaft orwith a helical turbine with a suction sleeve. This stirrer has athreefold function: to keep the solids in suspension, to draw in anddisperse the oxygen in the reaction mixture and to ensure, continuously,the recycling of the oxygen. The leaching tanks are also equipped withmeasuring and control devices which are not shown, for measuring thepotential within and the pressure above the reaction mixture and forregulating the oxygen flow rate as a function of the pressure and thestirrer speed as a function of the potential, or vice versa. These tanksare furthermore provided with a device, not shown, for checking thetemperature and with a safety valve.

Instead of being provided with a single multipurpose stirrer, theleaching tanks may be equipped with two stirrers: a constant-speedmixer-stirrer placed axially and used to keep the solids in suspensionand to disperse the oxygen, and a variable-speed self-sucking stirrerplaced eccentrically and used to recycle the unreacted oxygen. With thisarrangement, it would be advisable to regulate the oxygen flow rate as afunction of the potential and the speed of the self-sucking stirrer as afunction of the pressure.

The neutralization tank 35d is provided with a feed inlet, a spillway,means for regulating the flow rate of calcine as a function of theacidity and a device for checking the temperature.

In the plant described above, the first stage of leaching 29 isperformed in the first series of tanks and the second stage 30 in thesecond series of tanks.

The tank 33a is fed continuously with a stream of returning acid, withthe bottom stream 8 of a thickener, not shown, which separates theproducts of the neutral leaching 4, and with the solid phase 31 leavingthe filtration apparatus 36 which separates the products of the secondstage of leaching 30; the stream 8 therefore contains zinc ferrite andthe stream 32 the leaching residue which is rich in zinc, this residuealso containing zinc ferrite, especially the ferrite originating fromthe calcine used in the neutralization tank 35d.

The products of the first stage of leaching, which leave the tank 33d,are separated in the separator 34 and the stream 32 of primary leachatewhich is thus obtained is introduced continuously together with thestream 1b of zinc sulphite concentrate into the tank 35a.

The flow rates of the returning acid stream and of the streams 1b, 3band 8 are such that the molar ratio of the iron contained in the streams8 and 32 to the reactive sulphur contained in the stream 1b isapproximately 0.3 and that the sulphuric acid content of the streamleaving the tank 35c is approximately 20 g/l.

The pulp leaving the neutralization tank 35d has a sulphuric acidcontent of approximately 5 g/l.

The volumes of the tanks are such that the residence time of thereaction mixture is approximately 6 hours in the first series of tanksand approximately 5 hours in the second series of tanks.

In each leaching tank, the potential of the solution is maintained at anappropriate level, especially at 560-610 mV (SHE) in 33a, at 590-630 mVin 33b, at 610-650 mV in 33c, at 640-660 mV in 33d and at 560-620 mV in35a, 35b and 35c. The checking of the potential and, hence, thetrivalent iron content of the solution is performed by theabovementioned measuring and regulating devices.

The temperature in each leaching tank is kept at approximately 90° C.and the overpressure therein remains at a very low level, for example at5-20 cm of water, or even less, by virtue of the action of theself-sucking stirrer.

The action of the abovementioned measuring and regulating devices willnormally suffice to keep the potential at the intended level. However,if these devices were found for any reason whatsoever to be incapable bythemselves of keeping the potential at the intended level, it would alsobe possible to intervene by varying the temperature.

When working as described above, approximately 45% of the reactivesulphur is oxidized in the second stage of leaching and a zinc leachingyield of approximately 98% is reached, this being therefore with a totalleaching period of approximately 11 hours. The copper present in theconcentrate 1b is found again almost entirely in the leachate 19, fromwhich it will be subsequently separated, and the lead and the silverfrom the concentrate are found again in the leaching residue 25, fromwhich they can be easily separated by flotation, because this residue ispractically free from jarosites.

The streams 1a and 1b can obviously have the same composition or adifferent composition.

The number of tanks may vary. In fact, the leaching yield increases upto a certain point with the number of tanks, because with an increasingnumber of tanks it is possible to improve favourably the potentialprofile which it is desired to apply to the first stage of leaching andat the same time the probability that all the ore particles undergoleaching during the required period of time is increased. Needless tosay, however, the cost of the plant also increases with the number oftanks. The choice of this number will therefore be determined byconsiderations of a technical and economic nature.

A major advantage of the process of the invention, namely the shorteningof the duration of the operation (c), is illustrated by the examplesgiven below.

EXAMPLE 1

This example describes a test of leaching in one stage (operation (c))according to the process of the invention.

Starting materials employed

(α) 2 kg of a blende which has the following composition, in % byweight: 53.9 Zn, 5.6 Fe, 2.32 Pb, 30.5 S^(tot), 29.0 reactive S²⁻(=S^(tot) less pyrite S) and 0.02 Cu; this blende has a particle size of90% smaller than 44 mm;

(β) 1215 g of a zinc ferrite which has the following composition, in %by weight: 20.9 Zn, 30.4 Fe and 5.78 Pb;

(γ) 22.5 l of a cell returning acid containing 189 g/l of H₂ SO₄.

The molar ratio of the iron contained in (β) and the reactive sulphurcontained in (α) is 0.36.

Apparatus employed

A closed tank of 30-l capacity, equipped with a feed inlet, an oxygeninlet, a stirrer, a potentiometer probe and means for controlling thetemperature.

Leaching

(α) and (β) are added to (γ) over 60 minutes and at the same time thetemperature is gradually increased from 75° to 90° C. At the end of thisoperation virtually all of the ferrite has dissolved. Oxygen injectionis then commenced and leaching is continued. The reaction is stoppedafter 7.5 h.

Table 1 below gives the change in the main parameters during theleaching.

                  TABLE 1    ______________________________________    Time   mV         T     Fe.sup.2+                                    Fe.sup.3+                                         H.sub.2 SO.sub.4    h      (SHE)      °C.                            g/l     g/l  g/l    ______________________________________    1      590        90    10.6    0.2  120    2      593        90    14.0    0.6  89    3      595        90    15.4    0.7  79    4      597        90    15.6    0.8  70    5      603        90    16.1    1.0  65    6      610        90    15.4    1.6  60    7      617        90    14.8    2.1  56    7.5    625        90    14.4    2.6  53    ______________________________________

The pulp is filtered and 26.5 l of leachate which is rich in zinc and iniron and 1095 g of residue which is depleted in zinc are obtained.

The leachate which is rich in zinc and in iron contains, in g/l: 14.4Fe²⁺, 2.6 Fe³⁺ and 53 H₂ SO₄.

The residue which is depleted in zinc contains, in the dry state, in %by weight: 5.9 Zn, 1.3 Fe, 10.0 Pb, 57 S^(tot), 52 S° and 0.04 Cu.

The leaching yield of zinc is 95.2%.

EXAMPLE 2

This example describes a test of leaching in two stages (operation (c))according to the process of the invention.

Starting materials employed

(α) as in Example 1;

(β') 937 g of a zinc ferrite which has the same composition as that ofExample 1;

(γ') 14.8 l of a cell returning acid which has the same composition asthat of Example 1;

(δ) 1429 g of a leaching residue which is rich in zinc, which has thefollowing composition, in % by weight: 42.4 Zn, 4.5 Fe, 3.18 Pb, 42.9S^(tot), 21.5 reactive S²⁻, 18.8 S° and 0.05 Cu;

this residue was obtained during a previous operation which wassubstantially identical to the second stage of leaching which will bedescribed below, which means that 47.0% of the reactive sulphurcontained in (α) will be oxidized in this second stage of leaching.

The molar ratio of the iron contained in (β') to the reactive sulphurcontained in (α) is therefore 0.28, whereas the molar ratio of the ironcontained in (β') to the reactive sulphur contained in (δ) is 0.53.

Apparatus employed

As in Example 1, except that the closed tank has a capacity of 20 l.

First stage of leaching

First of all (δ) is added to (γ') over 30 minutes and then (β') over 60minutes while the temperature is gradually raised from 75° to 90° C.during the first hour of this charging operation. Oxygen is injectedduring the charging only when the potential of the pulp falls below 560mV. By first of all adding (δ') to (γ'), the potential of the solutionis lowered to a level of 560-610 mV, at which--as the Applicant hasascertained--zinc ferrite dissolves most quickly. (The cell returningacid (γ') has a potential appreciably higher than 610 mV. In a batchleaching, it is therefore important to take measures in order that thepotential of the acid should be rapidly returned to the level of 560-610mV. Such measures are generally not required in a continuous leachingbecause the pulp to which the cell returning acid, the zinc ferrite andthe leaching residue which is rich in zinc are added in this case willalmost always have a potential lower then 610 mV.)

Once the charging is finished, the introduction of oxygen into the tankis commenced and the potential of the solution is gradually raised byincreasing the flow rate of oxygen so as to obtain a value of 630-650 mvafter 6 h of leaching.

Table 2 below gives the change in the main parameters during this firststage of leaching.

                  TABLE 2    ______________________________________    Time   mV         T     Fe.sup.2+                                    Fe.sup.3+                                         H.sub.2 SO.sub.4    h      (SHE)      °C.                            g/l     g/l  g/l    ______________________________________    1      571        90    5       0.15 157    2      568        90    15.5    0.8  91    3      588        90    4      601        90    16.8    2.3  68    5      621        90    6      638        90    12.5    6.1  56    ______________________________________

The pulp is filtered and a primary leachate and 974 g of residuedepleted in zinc are obtained.

The primary leachate (ε) contains, in g/l: 12.5 Fe²⁺, 6.1 Fe³⁺ and 56 H₂SO₄.

The residue which is depleted in zinc contains, in the dry state, in %by weight: 3.2 Zn, 1.45 Fe, 9.2 Pb, 58 S^(tot), 55 S° and 0.03 Cu.

Second stage of leaching

The blende (α) is added continuously to the primary leachate (ε) over aperiod of time of 60 minutes while the temperature is raised at the sametime from 65° C. to 85° C. The oxygen flow rate is adjusted so as tokeep the potential of the solution between 560 and 590 mV. The leachingis stopped after 5 h.

Table 3 below gives the change in the main parameters during this secondstage of leaching:

                  TABLE 3    ______________________________________    Time   mV         T     Fe.sup.2+                                    Fe.sup.3+                                         H.sub.2 SO.sub.4    h      (SHE)      °C.                            g/l     g/l  g/l    ______________________________________    0.5               75    1      561        85    16.5    0.6  52    2      570        85    16.8    1.1  40    3      578        85    16.9    0.9  30    4      580        85    17.1    1.1  22    5      574        85    17.2    1.0  17    ______________________________________

After filtration of the pulp, a leachate which is rich in zinc and theleaching residue which is rich in zinc (δ) are obtained.

The leachate which is rich in zinc contains, in g/l: 17.2 Fe²⁺, 1.0 Fe³⁺and 17 H₂ S0₄.

The leaching yield of zinc is 98%, this being therefore after a leachingperiod of 11 hours.

EXAMPLE 3

This comparative example describes a test of leaching in two stages(operation (c)) according to the process of the prior art discussedabove (EP-A-0451456).

Starting materials employed

(α) as in Example 1;

(β") 1215 g of a zinc ferrite which has the same composition as that ofExample 1;

(γ") 16.6 l of a cell returning acid which has the same composition asthat of Example 1;

(δ') 1008 g of a leaching residue which is rich in zinc, which has thefollowing composition, in % by weight: 19.8 zn, 2.05 Fe, 4.5 Pb, 59S^(tot), 10.3 reactive S²⁻, 48 S° and 0.15 Cu; this residue was obtainedduring a previous operation which was appreciably identical to thesecond stage of leaching which will be described below, which means thatthis time 82.1% of the reactive sulphur contained in (α) will beoxidized in the second stage of leaching.

The molar ratio of the iron contained in (β") to the reactive sulphurcontained in (α) is here 0.36, that is to say a little higher andtherefore more favourable than in Example 2, whereas the molar ratio ofthe iron contained in (β") to the reactive sulphur contained in (δ') isnow 2.03.

Apparatus employed

As in Example 2.

First stage of leaching

The charging is performed as in Example 2, that is to say that first ofall (δ') is added to (γ") over 30 minutes and then (β") over 60 minuteswhile the temperature is gradually raised from 75° to 90° C. during thefirst hour. Leaching is then continued and is stopped after 4 h.

Attempts to lower the potential of the reaction mixture to the level of560-610 mV, which favours the dissolution of the ferrite, wereunsuccessful, probably because of the low content of reactive sulphur in(δ').

Table 4 below gives the change in the main parameters during this firststage of leaching.

                  TABLE 4    ______________________________________    Time   mV         T     Fe.sup.2+                                    Fe.sup.3+                                         H.sub.2 SO.sub.4    h      (SHE)      °C.                            g/l     g/l  g/l    ______________________________________    1      640        90    4.2     0.5  166    2      679        90    10.8    4.2  105    3      665        90    14.3    3.5  97    4      655        90    15.8    3.4  94    ______________________________________

The pulp is filtered and a primary leachate and 1079 g of residue whichis depleted in zinc are obtained.

The primary leachate (ε') contains, in g/l: 15.8 Fe²⁺, 3.4 Fe³⁺ and 94H₂ SO₄.

The residue which is depleted in zinc contains, in the dry state, in %by weight: 3 Zn, 1.7 Fe, 10.3 Pb, 56 S^(tot), 53 S° and 0.17 Cu.

Second stage of leaching

The blende (α) is added continuously to the primary leachate (ε') over aperiod of time of 60 minutes while at the same time the temperature israised from 65° C. to 85° C. The oxygen flow rate is adjusted so as tokeep the potential of the solution between 560 and 590 mV, as in Example2. However, after approximately nine hours' leaching, it is no longerpossible to keep the potential below 590 mV, which apparently means thatthe reactivity of the blende has become very low. Nevertheless, oxygencontinues to be injected in order to make the blende react further, andthe leaching is stopped after 16 h.

Table 5 below gives the change in the main parameters during this secondstage of leaching.

                  TABLE 5    ______________________________________    Time   mV         T     Fe.sup.2+                                    Fe.sup.3+                                         H.sub.2 SO.sub.4    h      (SHE)      °C.                            g/l     g/l  g/l    ______________________________________    1      567        85    19.70   0.1  101.5    2      579        85    20.30   0.35 92.5    4      589        85    21.20   0.60 75.5    5      589        85    22.20   0.90 57.0    10     598        85    22.95   0.90 52.2    12     611        85    22.40   1.90 41.5    14     613        85    22.35   2.45 32.5    16     619        85    21.85   3.25 22.0    ______________________________________

After filtration of the pulp, a leachate which is rich in zinc and theleaching residue which is rich in zinc (δ") are obtained.

The leachate which is rich in zinc contains, in g/l: 20.2 Fe²⁺, 2.8 Fe³⁺and 22 H₂ SO₄.

The leaching yield of zinc is 98%, this being therefore after a totalleaching period of 20 hours.

When these examples are compared, it is seen that the time required tocarry out the operation (c) in the process of the prior art exceeds by114% that required to carry out this operation with practically the sameyield in the first and second variants of the process of the invention,and by 82% that required to carry out this operation with the same yieldin the third and fourth variants of the process of the invention. Thisis equivalent to saying that, with the process of the invention, as muchis done in 0.47 reactor volume (1st and 2nd variants) or in 0.55 reactorvolume (3rd and 4th variants) as with the process of the prior art in 1reactor volume.

The industrial exploitation of the process of the invention willtherefore entail investment costs which will be far lower than those ofthe process of the prior art.

It is obvious that some special features of the operation (c) which havejust been described in connection with the process of the invention canbe very useful in a context other than that of the process of theinvention described above.

This is why the Applicant also requests protection for a process forleaching zinc ferrite together with a sulphide material containing zincsulphide, according to which the leaching is performed with a solutionof sulphuric acid at 60°-95° C. in atmospheric conditions so as toproduce a leachate which is laden with zinc and with iron and a leachingresidue which is depleted in zinc and in iron, this process beingcharacterized in that

(1) the work is done with a sulphide material:ferrite ratio such thatthe quantity of reactive sulphur present in the sulphide material isappreciably higher than that which can be oxidized by the iron presentin the ferrite, the reactive sulphur being the sulphur which is presentin the form of sulphide and which can be converted into elementalsulphur by the ferric sulphate,

(2) a stream of sulphide material, a stream of ferrite and a stream ofacid are introduced continuously into the first tank of a series oftanks, the pulp thus formed is passed successively through the othertanks of the series, a stream of oxygen is introduced into these othertanks and in each tank of the series conditions are maintained such thatthe pulp leaving the last tank consists of leachate laden with zinc andwith iron and of leaching residue which is depleted in zinc and in iron,and

(3) care is taken that the potential of the pulp should not fall below530 mV (SHE) and preferably not below 560 mV in the first tank.

The sulphide material may be a zinc sulphide concentrate or a partiallyleached zinc sulphide concentrate.

It is possible to refrain from introducing oxygen into the first tankand to keep the potential therein at at least 530 mv by working thereinwith a sulphide material:ferrite ratio which is sufficiently low and/orat a temperature which is sufficiently low.

It is also possible to keep the potential in the first tank at at least530 mv by introducing an appropriate stream of oxygen into it.

We claim:
 1. A process for leaching zinc ferrite together with asulphide material containing zinc sulphide, comprising the steps ofleaching the zinc ferrite and the sulfide material with a solution ofsulphuric acid at 60°-95° C. at a pressure equal to or differing by lessthan 20 kPa from atmospheric pressure to produce a leachate which isladen with zinc and with iron and a leaching residue which is depletedin zinc and in iron, quantities of the sulphide material and of zincferrite providing a molar ratio between iron contained in the zincferrite and reactive sulphur contained in the sulphide material to be atleast approximately 0.2, the reactive sulphur being the sulphur which ispresent in the form of sulphide and which can be converted intoelemental sulphur by oxidation with ferric sulphate, wherein,(1) theleaching is done with the sulphide material-ferrite ratio being suchthat the quantity of reactive sulphur present in the sulphide materialis higher than that which can be oxidized by the iron present in thezinc ferrite, (2) a stream of the sulphide material, a stream of thezinc ferrite and a stream of the sulfuric acid are introducedcontinuously into a first tank of a series of tanks forming a pulp inthe first tank which is passed successively through the other tanks ofthe series, a stream of oxygen is introduced into these other tanks andconditions are maintained in each tank of the series such that pulpleaving the last tank consists of leachate laden with zinc and withiron, containing 45-70 g/l of sulphuric acid and 2-10 g/l of Fe³⁺, andof leaching residue which is depleted in zinc and in iron, and (3) apotential of the pulp in the first tank is at least 530 mV (SHE). 2.Process according to claim 1, wherein the potential in the first tank iskept at at least 530 mV by controlling one of the sulphide material toferrite ratio and a tank temperature.
 3. Process according to claim 1,wherein the potential in the first tank is kept at at least 530 mV byintroducing a stream of oxygen into it.
 4. The process of claim 2wherein the potential is at least 560 mV in the first tank.
 5. In aprocess for extracting zinc from zinc sulphide concentrates comprisingthe steps ofa) providing zinc sulphide concentrates; b) roasting a firstportion of the concentrates to produce a calcine; c) neutral leaching afirst portion of the calcine to form a zinc rich and substantiallyiron-free leachate and a separate zinc ferrite residue; d) acid leachingthe separate zinc ferrite rich residue to form a zinc and iron richsolution; e) subjecting the zinc and iron rich solution to one of (1) areducing step to return a Fe⁺³ content of the zinc and iron richsolution to below 5 g/l by adding a second portion of the concentratesto the zinc and iron rich solution and a subsequent neutralizing stepwherein a remainder of the calcine is added to the zinc and iron richsolution to form a ferrite residue to be recycled to the acid leachingstep (d) and a solution containing iron and zinc with a free H₂ SO₄content below 10 g/l and (2) a neutralizing step wherein a remainder ofthe calcine is added to the zinc and iron rich solution to form aferrite residue to be recycled to the acid leaching step (d) and asolution containing iron and zinc with a free H₂ SO₄ content below 10g/l; f) precipitating iron from the solution containing iron and zincleaving a zinc rich solution and recycling the zinc rich solution to theneutral leach of step (c), the improvement comprising performing anadditional leaching step comprising one of:(i) leaching at least aportion of the step (c) separate zinc ferrite residue with a thirdportion of the concentrates, a return spent electrolyte at 60°-95° C. ata pressure equal to or differing by less than 20 kPa from atmosphericpressure and finely dispersed oxygen to produce a zinc and iron richleachate and a separate zinc-depleted leaching residue, the quantitiesof the third portion of the concentrates and the at least a portion ofthe step (c) separate zinc ferrite residue being such that a molar ratiobetween iron in the portion of the step (c) zinc ferrite residue and areactive sulfur contained in the third portion of the concentrates is atleast approximately 0.2, the reactive sulfur being sulfur which ispresent in the form of sulfide and which can be converted to elementalsulfur by oxidation with ferric sulphate; wherein leaching step (i) isperformed in a single stage so that the zinc and iron rich leachate hasa sulfuric acid content of 45-75 g/l and an Fe⁺³ content of 1-10 g/l andthe zinc and iron rich leachate is merged with one of the zinc and ironrich solution from the acid leaching step (d) and the zinc and iron richsolution of the reducing step (e)(1); and (ii) leaching at least aportion of the step (c) separate zinc ferrite residue with a thirdportion of the concentrates, a return spent electrolyte at 60°-95° C. ata pressure equal to or differing by less than 20 kPa from atmosphericpressure and finely dispersed oxygen to produce a zinc and iron richleachate and a separate zinc-depleted leaching residue, the quantitiesof the third portion of the concentrates and the at least a portion ofthe step (c) separate zinc ferrite residue being such that a molar ratiobetween iron in the portion of the step (c) zinc ferrite residue and areactive sulfur contained in the third portion of the concentrates is atleast approximately 0.2, the reactive sulfur being sulfur which ispresent in the form of a sulfide and which can be converted to elementalsulfur by oxidation with ferric sulphate; wherein the leaching step (ii)is performed in two stages;1) a first stage comprising leaching at leasta portion of the zinc ferrite residue and a recycled zinc rich leachingresidue with return spent electrolyte and finely dispersed oxygen toproduce a primary leachate and a zinc depleted residue; and 2) a secondstage wherein the third portion of the concentrates is leached with theprimary leachate and finely dispersed oxygen to produce a leachate richin zinc and iron and a separate zinc leaching residue, the zinc richresidue being recycled to the first stage, the second stage beingperformed in conditions such that less than 60% of the reactive sulphuris oxidized therein and that the zinc rich leaching residue producedtherein has a reactive sulphur content which is higher than that whichcan be oxidized in the first stage by the iron present in the zincferrite residue, the leachate rich in zinc and iron having a sulfuricacid content of 10-35 g/l and a Fe⁺³ content of 0.1-2 g/l, the leachaterich in zinc and iron being merged with one of the zinc and iron richsolution from the acid leaching step (d) and the zinc and iron richsolution of the reducing step (e)(1).
 6. The process of claim 5 whereinthe additional leaching step comprises step (i).
 7. The process of claim6 wherein all of the step (c) separate zinc ferrite residue is leachedin step (i) so that said acid leaching step (d) is bypassed.
 8. Theprocess of claim 6 wherein the sulphuric acid content is between 55 and65 g/l and the Fe⁺³ content is between 2 and 5 g/l for step (f)(i). 9.Process according to claim 6, wherein the zinc ferrite residue, thethird portion of the concentrate and the return spent electrolyte areintroduced continuously into a first tank of a series of tanks withoverflow which are placed in cascade, in each of these tanks conditionsare maintained such that pulp overflowing from the last tank consists ofthe zinc and iron rich leachate and of the zinc-depleted leachingresidue and the last tank overflows into a solid-liquid separator whichseparates the zinc-depleted leaching residue from the zinc and iron richleachate.
 10. The process of claim 5 wherein the additional leachingstep comprises step (ii).
 11. The process of claim 10 wherein all of thestep (c) separate zinc ferrite residue is leached in step (ii) so thatsaid acid leaching (d) is bypassed.
 12. The process of claim 10 whereinthe sulfuric acid content is between 10 and 25 g/l and the Fe⁺³ contentis between 0.5 and 1 g/l for step (f)(ii).
 13. Process according toclaim 10, wherein at least 30% of the reactive sulphur is oxidized inthe second stage of leaching.
 14. Process according to claim 13, whereinthe first stage of leaching is performed so that the Fe³⁺ concentrationof the solution reaches a value of 2-10 g/l in a final phase of thefirst stage.
 15. Process according to claim 14, wherein the Fe⁺³concentration is at least 0.1 g/l in an initial phase of the said firststage.
 16. Process according to claim 13, wherein the quantities of thereturn spent electrolyte, of the zinc ferrite residue and of reactivesulphur which are introduced into the first stage of leaching are suchthat the primary leachate has a sulphuric acid content of 40-70 g/l. 17.Process according to claim 13, characterized in that the Fe³⁺concentration of the solution in the second stage of leaching ismaintained constantly at a level of 0.1-2 g/l.
 18. Process according toclaim 10, wherein prior to separating the leachate which is rich in zincand in iron from the leaching residue which is rich in zinc, theremainder of the calcine is added during a final chase of the leachingstep for adjustment of the free H₂ SO₄ content so that step (e) is bypassed.
 19. Process according to claim 18, wherein, during the firststage leaching, the zinc ferrite residue, the leaching residue which isrich in zinc and the return spent electrolyte are introducedcontinuously into a first tank of a first series of tanks with overflowwhich are placed in cascade, in each of these tanks conditions aremaintained such that pulp overflowing from the last tank comprises theprimary leachate and the zinc-depleted residue, the last tankoverflowing into a solid-liquid separator which separates thezinc-depleted residue from the primary leachate, this primary leachatetogether with the third portion of the concentrates are introducedcontinuously into a first tank of a second series of tanks with overflowwhich are placed in cascade, in each tank of this second seriesconditions are maintained such that pulp overflowing from the last tankconsists of the leachate rich in zinc and in iron and leaching residuerich in zinc, the last tank of this series overflowing into aneutralization tank in which the remainder of the calcine is added tothe pulp, the neutralization tank overflowing into a solid-liquidseparator which separates the leachate which is rich in zinc and ironfrom the leaching residue rich in zinc, the leaching residue then beingrecycled towards the first tank of the first series.
 20. Processaccording to claim 9, wherein the tanks are equipped with a self-suckingstirrer, are connected to a source of oxygen, and a stirrer speed andoxygen delivery are regulated so as to keep a potential in each tank ata set level.
 21. The process of claim 5 wherein the Fe⁺³ content isbetween 1 and 3 g/l for said reducing step and the free H₂ SO₄ contentis between 3 and 5 g/l for said neutralizing step.
 22. The process ofclaim 13 wherein at least 40% of the reactive sulfur is oxidized in thesecond stage of leaching.
 23. The process of claim 14 wherein the Fe⁺³concentration is between 3 and 7 g/l in the final phase.
 24. The processof claim 15 wherein the Fe⁺³ concentration is at least 0.2 g/l in theinitial phase.
 25. The process of claim 17 wherein the Fe⁺³concentration in the second stage is at a level between 0.5 and 1 g/l.26. The process according to claim 19, wherein the tanks of the firstand second series are equipped with a self-sucking stirrer, areconnected to a source of oxygen, and a stirrer speed and oxygen deliveryare regulated so as to keep a potential in each tank at a set level.